Process for the manufacture of electrolytic copper

ABSTRACT

A process is described for making electrolytic copper of high quality from chalcopyrite flotation concentrates, and the like. The concentrates are first slurried and then subjected to a high temperature-high pressure oxidation-leaching operation in an autoclave system under controlled conditions. The resulting slurry from the oxidation-leaching operation is flashed to atmospheric pressure in order to remove heat and then subjected to a solids-liquid separation. The separated liquor from the solids-liquid separation system, containing the recovered copper values, is cooled and then fed to an electrolytic deposition operation where metallic copper cathodes of high purity are produced. The spent electrolyte is further subjected to a sulfide treatment to precipitate and recover, as copper sulfide, the residual amounts of copper not deposited in the electrolytic cells. The copper sulfide is then recycled in the form of a slurry to the initial stages of the process.

United States Patent [1 1 Fisher et a].

Nov. 4, 1975 PROCESS FOR THE MANUFACTURE OF Primary Examiner-R L.Andrews ELECTROLYTIC COPPER Attorney, Agent, or Firm-Lawrence W. Flynn[75] Inventors: Bernard M. Fisher, New Orleans;

Robert C. Hills, Metairie; Freddie J. Touro, New Orleans, all of La.[57] ABS CT P A process is described for making electrolytic copper [73]Asslgnee' fi f g rals Company New of high quality from chalcopyriteflotation concen- Or trates, and the like. The concentrates are firstslurried [22] Filed: Feb. 27, 1974 and then subjected to a hightemperature-high pressure oxidation-leaching operation in an autoclavesys- [21] Appl. No.: 446 314 tem under controlled conditions, Theresulting slurry from the oxidation-leaching operation is flashed to at-[52] U.S. Cl. 204/108; 75/1 17; 423/140; mosph ric pressur in order oremove hea and then 423/633; 423/602; 423/158; 423/37; 423/87; subjectedto a solids-liquid separation. The separated 423/522; 423/555; 423/35;423/41 liquor from the solids-liquid separation system, con- [51] Int.Cl. C22d 1/16; C22b 15/00 taining the recovered copper a is Cooled and[58] Field of Search 75/1 17; 423/140, 158, 633, then fed to anelectrolytic deposition operation where 423/602, 37, 87, 522, 555, 41;204/108 metallic copper cathodes of high purity are produced. The s entelectrol te is further sub'ected to a sulfide P Y J [56] ReferencesCited treatment to precipitate and recover, as copper sul- UNITED STATESPATENTS fide, the residual amounts of copper not deposited in theelectrolytic cells. The copper sulfide is then recy- 3,637,371 1/1972Maclaw et al. 75/117 ded in the form 0f a slurry to the initial Stagesof the process.

22 Claims, 1 Drawing Figure OXYGEN NEUTQALIZEH J 4 CONCENTHAYE SLURRY lOX DATlON- TANK LEACHING SLURRVED CONCENTRATE 2 mass, TREATED $LURRY 5COO AC'OlC SLURQY new, a

STEAM =ifig gg= E F LASHI NG 4 ermis f f wn'znll b PLUSGKIMN l COOLINGCOOLED LIQUOR l2 SEFARATED 1 -s|.unflv 9 uqvon l ELECTROLVTlCSOLIDS-LIQUID DEPOSITION com, SEPARATION PRDOUCT re SPENT lELEC]T9F\OLYTE l :SULFIDING AGENT 2o *SULFIDED SLURRV FLOW WASHINGSULFIDE k s w m MT TH'CKENER l 124 ;112:1195 15 LVMEESAONE sY suM SLURRV(WASTE) U.S. Patent Nov. 4, 1975 3,917,519

OXY3GEN NEUTRALIZER ORE 4 CONCENTRATE {1L1 II I E SLURRY OXIDATlON- TANKI LEACHING SLURRIED CONCEgTRATE COOLING WATER OUT TREATED 6 SLUSRRY f!COOLING AcIOIc SLURRY f OF COPPER COOUNG a SULFlDE WATER nq 4 SLURRY 722 6 F I STEAM 8 E ZZ L E F LASH I NG I (WASH WAZER S Y COPPER v LUE IWATER H PLUSGACID) 1 I COOLING COOLED LIQUOR l2 SEPARATED f lswamrgLIQUOR ELECTROLYTIC SOLIDS-LIQUID DEPOSITION COPPER SEPARATION PRODUCT18 SPENT I --EL.ECTROLYTE suu-Ioms AGENT 2o SULFIDED SLURRY v OVERFLOWWASHING SULF'DE c lprim ms WATER THICKENER I IMPURITIES 15 23 I *1 I f ILIMESTONE WASHED l 24 SOLIDS 17 GYPSUM SLURRY WASTE) 25 PROCESS FOR THEMANUFACTURE OF ELECTROLYTIC COPPER BACKGROUND OF THE INVENTIONincreasingly, in recent years, consideration has been given to employinghydrometallurgical processes for the recovery of metal values fromsulfide ores. One of the reasons for this is that hydrometallurgicalprocesses, for the most part, do not involve the generation of sulfurdioxide and the consequent problems of air pollution characteristic ofpyrometallurgical process. Another reason for consideringhydrometallurgical processes is the extreme flexibility afforded withregard to the virtually infinite sets of unique conditions oftemperature, pressure, retention time, specific solvents and additivesand procedures that can be resorted to for whatever particularmetallurgical separation one is confronted with.

Where the final step in a hydrometallurgical copper recovery process isto be based upon electrowinning, it would normally be desirable that thepreliminary steps for producing the electrolyte include the bestpossible solids-liquid separation, reduction ofdissolved iron andarsenic to minimal levels, and concentration of copper, as coppersulfate, to maximum levels in the feed to the electrowinning step. Whilea variety of methods are available for carrying out each of these steps,in most known processes these individual methods do not alwayscomplement each other, i.e., they lack in technological compatibilitywith regard to making up an efficient, economical, pollution-freeprocess. It is an object of this invention to provide a process whereineach individual step is uniquely suited to the others and to the overallprocess for manufacturing a high quality metallic copper productefficiently and economically. A sec ondary object is to circumvent theproduction of sulfur dioxide and thereby negate the possiblity of airpollution.

BRIEF SUMMARY OF THE INVENTION In accordance with and furtherance ofthis stated objective the present invention provides a process wherein acopper sulfide flotation concentrate, at least a portion of which ischalcopyrite, CuFeS is first mixed with water and recycle processliquors and slurries. Alternatively, the concentrate may be slurriedwith recycle process liquors and slurries only. The resulting slurry isthen subjected to a high temperaturehigh pressure operation in anautoclave system, wherein the temperature is closely controlled and anoxygen partial pressure (from industrial oxygen or an oxygen bearinggas) is maintained. The autoclave sys tern may be a single vessel or aseries of vessels The oxidation in the high temperature-high pressureoperation results in the formation of sulfuric acid and thesolubilization of the metal constituents of the concentrate in the formof sulfates. In addition to copper, iron and arsenic also are dissolvedby the acid, but the latter two are removed from solution, aftersolubilization of virtually all of the copper, by the neutralization ofthe greater portion of the sulfuric acid with a neutralizing agent suchas limestone. The neutralization of the sol furic acid results in theprecipitation of virtually all of the dissolved iron and arsenic.

Exiting the high temperature-high pressure opera tion, the slurry isflashed to atmospheric pressure and fed to a solids-liquid separationsystem. The separated liquor from the solids-liquid separation system isthen cooled and fed to an electrolytic deposition operation whereend-product copper cathodes are produced. The solids from thesolids-liquid separation operation are normally washed with water andthe return wash water, containing residual copper values, is recycledback to the initial stages of the process. The washed solids, composedof the ore gangue and the materials precipitated in the autoclave-Fe O(hematite), iron and calcium arsenates, and CaSO (anhydrite)are sent onfor further processing or to waste. The spent electrolyte from theelectrolytic deposition operation is then treated with a sulfidingagent, such as hydrogen sulfide, to precipitate, as copper sulfide, thecopper values not deposited with the copper cathodes. The copper sulfideslurry is then thickened to a solids content of between about 3 and l5percent and recycled back to the initial stages of the process.

It is significant of the process of this invention that the varioussteps required are uniquely compatible with each other and with theoverall process. For example, the high-temperature precipitation of ironand calcium as hematite and anhydrite, respectively, makes thesubsequent solids-liquid separation much easier than if the precipitatedmaterials were ferric hydroxide and gypsum as would be obtained in alow-temperature processv Additionally, the steps wherein water vapor isremoved from the liquor in the process of venting noncondensable gases,or controlling the temperature, or flashing, etc., are compatible withthe overall object of the leaching step, that is, providing a liquorwith a maximum concentration of copper as copper sulfite and a very lowconcentration of dissolved iron.

These and other aspects of the invention will be understood morethoroughly in the light of the following description, as illustrated inthe accopanying drawing.

BRIEF DESCRIPTION OF THE DRAWING The drawing is a flow sheetillustrating the preferred embodiment of the invention.

DETAILED DESCRIPTION The preferred embodiment of the invention involvesthe treatment of a chalcopyrite flotation concentrate and is bestillustrated with reference to the drawing. Referring to the drawing,than, a chalcopyrite flotation concentrate 1, containing between about20 and 30% copper, is fed into a slurry tank and mixed with recycle washwater 16 containing residual amounts of copper sulfate and sulfuricacid, and with a recycle acidic slurry of copper sulfide 22. Theresultant slurry 2 is fed continuously to a high temperature-highpressure oxidation-leaching operation which, in a preferred embodiment,is carried out in a multi-compartment, horizontal autoclave systemequipped with agitators.

Withn the autoclave system the slurry temperature is maintained between350 and 450F, and preferably between 425 and 450F. The total pressure ismaintained at between about 300 and I000 psig, and preferably betweenabout 400 and 600 psig, by the introduction of a high qualityoxygen-containing gas 3 so as to provide an oxygen partial pressure ofbetween about 50 and 500 psi, and preferably between about 100 and 200psi. The oxygencontaining gas used should have an oxygen content of atleast about and preferably in the order of 99%.

The oxidizing chemical reactions in the autoclave system result in theformation of sulfuric acid and the dissolution of the copper and iron assulfates. Because the oxidizing reactions are exothermic, heat must beremoved to maintain the temperature within the desired range. Water maybe removed from the system during the oxidation-leaching operation,resulting in a desirable increase in the concentration of the sulfatesin the slurry liquor. The autoclave system may be a single pressurizedvessel or may include several pressurized vessels.

A suitable process and apparatus for the removal of heat and waterduring the oxidation-leaching operation is described in copendingapplication for US. Pat. Ser. No. 446,412, filed Feb. 27, 1974 byFreddie J. Touro, one of the present inventors.

Removal of water also serves to increase the concentration of theacidity of the slurry, which is desirable, since increased acidityresults in increased copper extraction rates when the oxidation-leachingoperation is carried out under the preferred temperature conditions of425450F. Control of the acidity during the oxidation-leaching ofchalcopyrite forms the subject matter of another copending applicationfor US. Pat. Ser. No. 446,315, filed Feb. 27, 1974 by Freddie .I. Touro,one of the present inventors.

To insure maximum removal of iron from the leach liquor, a neutralizingagent such as limestone slurry 4 is added in the autoclave system so asto reduce the acidity to less than about 40 grams per liter H 80preferably between about IO and 20 grams per liter H 80 and permit thehydrolysis of Fe (SO and precipitation of iron as Fe O- Neutralizationof the acidity at a temperature between 425450F in the autoclave is aparticularly preferred embodiment of the process because in thistemperature range the iron is precipitated as hematite (Fe O and thecalcium as anhydrite (CaSO both of which are crystalline in nature andtheir precipitation in this form results in the satisfactory separationof the solids from the liquid later in the process. In contrast, theprecipitates obtained at low temperatures (iron as ferric hydroxide, andcalcium as gypsum) are much more difficult to separate from the liquidphase.

While limestone, i.e., calcium carbonate, has been described as theneutralizing agent used in the preferred embodiment, other neutralizingagents may be utilized. For example, the hydroxides, oxides andcarbonates of calcium, strontium and barium, all of which form insolublesulfate, may be used as the neutralizing agent.

Following the addition of the neutralizing agent, the slurry is providedadditional retention tim in the autoclave system at the preferredtemperature of between 425 and 425F, and at the stated oxygen partialpressure with constant mechanical agitation. During this time,additional iron is precipitated and further oxidation and dissolution ofcopper may occur. A vaporspace bleed-stream may be vented to atmospherefrom the autoclave to purge the gradual buildup of contaminant nitrogen(introduced with the oxygencontaining gas). Some oxygen is lost withthis bleedstream as well as some water vapor. The loss of the latteragain serves to increase the copper concentration in the leach liquorand to remove some heat.

In the preferred embodiment the treated slurry stream 5 leaving theautoclave is made up of a liquid phase composed of a solution of acidiccopper sulfate and solids composed of gangue, anhydrite (CaSO andhematite (Fe O Before this liquid phase is separated from these solids,the slurry is subjected to a flashing operation. The flashing operationmay optionally be preceded by an indirect cooling operation. In thepreferred embodiment, the flashing operation is preceded by an indirectcooling operation, as indicated in the drawing. Exiting the autoclave,then, the slurry is first cooled with water 6 in an indirect heatexchanger to approximately 275400F while maintaining the same pressureas used in the autoclave. The heat exchanger may be a waste heat boiler,for example, generating 15 psig steam. This reduction in temperature isprovided so that when the cooled slurry 7 is finally flashed toatmospheric pressure, the volume of flashed steam produced 8 issignificantly less than that which would be produced if flashingoccurred at the original temperature of between 425 and 450F. As aresult, the velocity of the slurry -steam mixture through the letdownvalve is decreased and the erosion of the valve reduced in severity.

In the process of being flashed to atmospheric pressure, additionalwater vapor is lost and the copper concentration of the liquor furtherincreased. After flashing, the temperature of the slurry isapproximately 230F (somewhat elevated for atmospheric pressure due tothe content of dissolved salts). Further cooling of the liquor, down toapproximately 150F, may be carried out before the separation of theliquid and solids by blending recycle liquor l3, cooled to about 120F ina subsequent cooling operation (described below), with the 230F slurry.The resultant slurry 9, cooled to about 150F, is then fed to asolids-liquid separation system with separated liquor 10 going to acooling operation and solids 14 going to a washing system. In apreferred embodiment a barometric condenser is used to cool stream 10 toabout 120F. A portion of the cooled separated liquor 13 is recycled tocool the slurry, as it leaves the flashing operation, as describedabove. In the barometric condenser additional water 11 is removed undervacuum at about 120F to further increase the copper concentration of theleach liquor. Removal of water at this point has the beneficial effectof increasing the wash water-to-tailings ratio that may be used in thewashing system.

In the washing system the solids are washed with water 15 to recover theretained copper sulfate solution. The return wash water 16, containingcopper values and dilute sulfuric acid, is recycled back to the slurrypreparation facilities. The washed solids tailings 17 may be sent to atailings pond or to further treatment for the recovery of gold orsilver, etc. It is to be noted that any type of solids-liquid separationtechniques, such as filtration, centrifuging, etc., may be employed inthese operations; however, a thickener or a combination of thickeners ispreferred. In the preferred embodiment, the portion of the I20F liquorfrom the barometric condenser that is not recycled 12 is fed to theelectrolytic cells. At this point, the concentration of copper in theseparated copper sulfate solution has been increased to about grams perliter.

In the electrolytic deposition process copper cathodes 18 of as high as99.9 percent purity can be produced as final product. Heat is liberatedduring this or eration and should be removed in order to keep thetemperature in the electrolytic cells at about 150F or less. Sulfuricacid is also produced to the stoichiometric extent of the copperdeposited so that the spent electrolyte l9 normally contains betweenabout 5 and 15 grams per liter of copper and between about and 170 gramsper liter of sulfuric acid.

The spent solution 19 is then treated with a sulfiding agent 20 toprecipitate the copper as copper sulfide. Any one of a number ofsulfiding agents which cause the precipitation of the copper from spentelectrolyte 19 as copper sulfide, such as hydrogen sulfide (H 8),ammonium sulfide (NH S), sodium sulfide (Na S), ammonium hydrosulfide(NH HS), sodium hydrosulfide (NaI-IS), potassium sulfide (K 5) andpotassium hydrosulfide (Kl-IS), may be used. Preferably hydrogen sulfide(H 8) is used. Preferably an amount in excess of that stoichiometricallyrequired is employed in order to assure that precipitation of virtuallyall of the copper present in the spent electrolyte. The sulfided slurry21 is processed in a thickener and the thickened acidic copper sulfide(CuS) slurry 22, having a solids content of between about 3 percent andpercent and preferably higher than 5 percent, is recycled to the feedslurry tank. About 70-90% of spent electrolyte stream 19 can be bledfrom the process via stream 23 by this method with minimal loss ofcopper. This bleed serves to eliminate from the process certain solubleimpurities such as calcium, magnesium, nickel and cobalt. The acidicsupernatant liquor stream 23 from the thickener may be treated with alimestone slurry 24, as shown in the drawing, and the resultant gypsumslurry 25 pumped to waste, i.e., to holding ponds or land-fill areas, orit may be used in any other process which requires the use of a dilutesulfuric acid solution.

Example The following example illustrates the manner in which theprocess of this invention may be operated.

Chalcopyrite concentrate 1 may be fed to the system of the drawing andmixed with recycle streams l6 and 22 as previously indicated. Thecomposition of stream 1 is given below, in Table 1, together with thetypical compositions of other pertinent streams of the process.

Oxidation-leaching of the slurried concentrate 2 is carried out at 425Fand 100 psi of oxygen partial pressure in an autoclave system. Totalpressure is 445 psig. Limestone 4 is used to partially neutralize theacidity during oxidation-leaching so that the exiting slurry 5 goin Jthe indirect cooler has an acidity of about grams per liter H 80 Thetemperature of slurry 5 is decreased in the indirect cooler to about320F. Partially cooled slurry 7 is then flashed to atmospheric pressureand to a temperature of about 230F, and blended with cooled recycleliquor stream 13 for further cooling to 150F before entering thesolids-liquid separation system.

The solids-liquid separation is carried out in one thickener, and thebottoms 14 from this thickener are washed with water in a countercurrentfashion using three thickeners. Tailings 17 from the last washing 6thickener may be sent to waste. The overflow 10 from the solids-liquidseparation thickener is further cooled to 120F using a barometriccondenser and then divided into stream 13, which is used for blendingwith the flashed slurry, as described above, and stream 12, which is fedto the electrolytic deposition operation.

Copper cathodes 18 of high purity are produced in the electrolyticcells. The temperature in this operation normally tends to rise and soan external cooler (not shown) is provided to maintain it at about I50F.In the electrolytic deposition operation the copper content of theelectrolyte is reduced from about -80 grams per liter (in stream 12) toabout 5 grams per liter (in stream 19). Depletion of the copper, duringelectrolysis, to below about 5 grams per liter is not deemedeconomically attractive in view of the convenience of scavenging thecopper values not deposited during electrolysis which is provided by thesubsequent sulfiding operation of our process.

Spent electrolyte 19 is treated with hydrogen sulfide 20 in an amountsufficient to precipitate as copper sulfide virtually all of the copperpresent in solution at this point. Sulfided slurry 21 is thickened andrecycled as stream 22 to the slurry feed tank. The overflow stream 23from the sulfide thickener contains most of the impurities not rejectedin the oxidation-leaching as precipitated solids and is convenientlytreated with limestone 24 and pumped to waste 25. It is significant ofthe process of this invention that these impurities are rejected fromthe system in the manner just described and without losing anysignificant amount of copper values. By contrast, a process in which thespent electrolyte 19 is sent to waste would lose significant amounts ofcopper values as dissolved copper in the waste spent electrolyte.

It is also significant of the process of this invention that the coppervalues not deposited in the electrolytic deposition operation areconveniently recovered by the recycling of a relatively concentratedcopper sulfide slurry 22 to the initial stages of the process withoutthe recycling of impurities or large quantities of undesirable water tothe initial stages of the process. By contrast, a process which chose torecycle the spent electrolyte 19 from the electrolyte depositionoperation in order to recover the copper values not deposited duringelectrolysis would do so at the expense of recycling the impuritiesassociated with the spent electrolyte and, of course, of not being ableto remove water from the system at this point. The inability to re movewater from the system at this point would have the effect, in such case,of decreasing the wash water-to-tailings ratio which may be used in thewashing operation of the process and this, in turn, would increase thenumber of thickeners necessary to effectively carry out the washing.

TABLE 1 STREAM COMPOSITIONS OF A TYPICAL RUN OF PROCESS FOR THEMANUFACTURE OF ELECTROLYTIC COPPER TABLE l-continued STREAM COMPOSITIONSOF A TYPlCAL RUN OF PROCESS FOR THE MANUFACTURE OF ELECTROLYTIC COPPERAll stream numbers refer to the drawing. "Compositions of streams No. I,I7 and 22 are given on a dry basis.

The terms and expressions which have been used here are used as terms ofdescription and not of limitation, and there is no intention, in the useof such terms and expressions, of excluding equivalents of the featuresshown and described, or portions thereof, it being recognized thatvarious modifications are possible within the scope of the inventionclaimed.

What is claimed is:

1. Process for recovering copper from a copper-bearing sulfide orecontaining iron comprising the steps of:

passing an aqueous slurry of said ore through an oxidation-leachingoperation in an autoclave system while maintaining the temperaturewithin said autoclave system at between 350F and 450F and providing anoxygen partial pressure within said autoclave system of between about 50and 500 pounds per square inch to form sulfuric acid and to solubilizethe metal constituents in said slurry as sulfate solutions,

reducing the acidity of said solutions in said autoclave system by theaddition of a neutralizing agent during said oxidation-leachingoperation to precipitate substantially all of the solubilized iron insaid solutions as insoluble ferric oxide,

withdrawing from said autoclave system the treated slurry comprising aliquid phase of acidic copper sulfate and solids comprising ore gangueand the precipitated ferric oxide,

flashing the treated slurry to atmospheric pressure,

separating said liquid phase from said solids,

washing said separated solids with an aqueous wash medium to recoverretained copper sulfate solution and recycling said recovered coppersulfate solution to said autoclave system,

cooling said separated liquid phase,

electrolytically treating said separated liquid phase to produceelectrolytic copper and a spent electrolyte comprising sulfuric acid anda residual amount of copper,

treating said spent electrolyte with a sulfiding agent to precipitatesubstantially all of said residual amount of copper from saidelectrolyte as copper sulfide,

thickening said treated spent electrolyte to separate said precipitatedcopper sulfide from said treated spent electrolyte as a thickened coppersulfide slurry, and

recycling said thickened copper sulfide slurry to said autoclave system.

2. The process as defined in claim 1 wherein water vapor produced duringsaid oxidation-leaching operation within said autoclave system isremoved from said autoclave system whereby the concentration of saidsulfate solutions is increased.

3. The process as defined in claim 1 wherein the acidity of saidsolutions in said autoclave system is reduced to less than 40 grams perliter of sulfuric acid by the addition of said neutralizing agent.

4. The process as defined in claim 1 further comprising the step ofcooling the treated slurry withdrawn from said autoclave system tobetween about 275F and 400F while maintaining the same total pressureused in said oxidation-leaching operation prior to said flashingoperation.

5. The process as defined in claim 1 further comprising the step ofcooling the treated slurry to approximately l50F prior to saidseparating step.

6. The process as defined in claim 5 wherein the concentration of copperin said separated liquid phase following said cooling and water removalstep is about grams per liter of copper.

7. The process as defined in claim 1 wherein water is removed from saidseparated liquid phase during said cooling step.

8. The process as defined in claim 1 wherein a portion of said cooledseparated liquid phase is blended with said treated slurry prior toseparating said liquid phase from said solids.

9. The process as defined in claim 1 wherein the temperature of saidseparated liquid phase is reduced to about 120F by said cooling step.

10. The process as defined in claim 1 wherein said cooling of saidseparated liquid phase is carried out using a barometric condenser.

11. The process as defined in claim 1 wherein said spent electrolyte hasa concentration of between about 5 and 15 grams per liter of copper andbetween about 100 and I70 grams per liter of sulfuric acid.

12. The process as defined in claim 1 wherein said sulfiding agent ishydrogen sulfide.

13. The process as defined in claim 1 wherein said sulfiding agent isselected from the group consisting of hydrogen sulfide, ammoniumsulfide, sodium sulfide, ammonium hydrosulfide, sodium hydrosulfide,potassium sulfide, and potassium hydrosulfide.

14. The process as defined in claim 1 wherein said treated spentelectrolyte is thickened to a solids content of between about 3 and 15percent.

15. The process as defined in claim 1 wherein the copper content of saidliquid phase is reduced from between about 60 and grams per liter toabout 5 grams per liter in said electrolytic treatment.

16. The process as defined in claim 1 wherein the acidity of saidsolutions in said autoclave system is reduced to between about 10 and 20grams per liter of sulfuric acid by the addition of said neutralizingagent.

17. The process as defined in claim 1 wherein the oxidation-leachingoperation is carried out at a temperature between 425F and 450F.

18. The process as defined in claim 1 wherein the oxi dation-leachingoperation is carried out using an oxygen partial pressure of betweenabout I00 and 200 wherein a first part of said calcium in said calciumcarbonate forms insoluble calcium sulfate, wherein said arsenic,solubilized to a metal sulfate in said oxidationleaching operation, isprecipitated with at least a part of said calcium as insoluble iron andcalcium arsenates upon the addition of said calcium carbonate, andwherein said iron and calcium arsenates are withdrawn from saidautoclave as part of the solids in said treated slurry.

22. The process as defined in claim 1 wherein the electrolytic copperproduced is about 99.9 percent pure.

1. PROCESS FOR RECOVERING COPPER FROM A COPPER-BEARING SULFIDE ORECONTAINING IRON COMPRSING THE STEPS OF: PASSING AN AQUEOUS SLURRY OFSAID ORE THROUGH AN OXIDATIONLEACHING OPERATION IN AN AUTOCLAVE SYSTEMWHILE MAINTAINING THE TEMPERATURE WITHIN SAID AUTOCLAVE SYSTEM ATBETWEEN 350*F AND 450*F AND PROVIDING AN OXYGEN PARTIAL PRESSURE WITHINSAID AUTOCLAVE SYSTEM OF BETWEEN ABOUT 50 AND 500 POUNDS PER SQUARE INCHTO FORM SULFURIC ACID AND TO SOLUBILIZE THE METAL CONSTITUENTS IN SAIDSLURRY AS SULFATE SOLUTIONS. REDUCING THE ACIDITY OF SAID SOLUTIONS INSAID AUTOCLAVE SYSTEM BY TH ADDITION OF A NEUTRALIZING AGENT DURING SAIDOXIDATION-LEACHING OPERATION TO PRECIPITATE SUBSTANTIALLY ALL OF THESOLUBILIZED IRON IN SAID SOLUTIONS ASINSOLUBLE FERRIC OXIDE, WITHDRAWINGFROM SAID SUTOCLAVE SYSTEM THE TREATED SLURRY COMPRISING ALIQUID PHASEOF ACIDIC COPPER SULFATE AND SOLIDS COMPRISING ORE GANGUE AND THEPRECIPITATED FERRIC OXIDE, FLASHING THE TREATED SLURRY TO ATMOSPHERICPRESSURE, SEPARATING SAID LIQUID PHASE FROM SAID SOLIDS, WASHING SAIDSEPARATED SOLIDS WITH AN AQUEOUS WASH MEDIUM TO RECOVER RETAINED COPPERSULFATE SOLUTION AN RECYCLING SAID RECOVERED COPPE SULFATE SOLUTION TOSAID AUTOCLAVE SYSTEM, COOLING SAID SEPARATED LIQUIDD PHASE,ELECTROLYTICALLY TREATING SAID SEPARATE LIQUID PHASE TO PRODUCEELECTROLYTIC COPPER AND A SPENT ELECTROLYTE COMPRISING SULFURIC ACIDDDADA RESIDUAL AMOUNT OF COPPER, TREATING SAID SPENT ELECTROLYTE WITH ASULFIDING AGENT TO PRECIPITATE SUBSTANTIALY ALL OF SAID RESIDUAL AMOUNTOF COPPER FROM SAID ELECTROLYTEAS COPPER SULFIDE, THICKENING SAIDTREATED SPENT ELCTROLYTE TO SEPARATE SAID PRECIPITATED COPPER SULFIDEFROM SAID TREATEDSPENT ELECTROLYTE AS A THICKENED COPPE SULFIDE SLURRY,AND RECYCLING SAID THICKENED COPPE SULFIDE SLURRY TO SAID AUTOCLAVESYSTEM.
 2. The process as defined in claim 1 wherein water vaporproduced during said oxidation-leaching operation within said autoclavesystem is removed from said autoclave system whereby the concentrationof said sulfate solutions is increased.
 3. The process as defined inclaim 1 wherein the acidity of said solutions in said autoclave systemis reduced to less than 40 grams per liter of sulfuric acid by theaddition of said neutralizing agent.
 4. The process as defined in claim1 further comprising the step of cooling the treated slurry withdrawnfrom said autoclave system to between about 275*F and 400*F whilemaintaining the same total pressure used in said oxidation-leachingoperation prior to said flashing operation.
 5. The process as defined inclaim 1 further comprising the step of cooling the treated slurry toapproximately 150*F prior to said separating step.
 6. The process asdefined in claim 5 wherein the concentration of copper in said separatedliquid phase following said cooling and water removal step is about 75grams per liter of copper.
 7. The process as defined in claim 1 whereinwater is removed from said separated liquid phase during said coolingstep.
 8. The process as defined in claim 1 wherein a portion of saidcooled separated liquid phase is blended with said treated slurry priorto separating said liquid phase from said solids.
 9. The process asdefined in claim 1 wherein the temperature of said separated liquidphase is reduced to about 120*F by said cooling step.
 10. The process asdefined in claim 1 wherein said cooling of said separated liquid phaseis carried out using a barometric condenser.
 11. The process as definedin claim 1 wherein said spent electrolyte has a concentration of betweenabout 5 and 15 grams per liter of copper and between about 100 and 170grams per liter of sulfuric acid.
 12. The process as defined in claim 1wherein said sulfiding agent is hydrogen sulfide.
 13. The process asdefined in claim 1 wherein said sulfiding agent is selected from thegroup consisting of hydrogen sulfide, ammonium sulfide, sodium sulfide,ammonium hydrosulfide, sodium hydrosulfide, potassium sulfide, andpotassium hydrosulfide.
 14. The process as defined in claim 1 whereinsaid treated spent electrolyte is thickened to a solids content ofbetween about 3 and 15 percent.
 15. The process as defined in claim 1wherein the copper content of said liquid phase is reduced from betweenabout 60 and 80 grams per liter to about 5 grams per liter in saidelectrolytic treatment.
 16. The process as defined in claim 1 whereinthe acidity of said solutions in said autoclave system is reducEd tobetween about 10 and 20 grams per liter of sulfuric acid by the additionof said neutralizing agent.
 17. The process as defined in claim 1wherein the oxidation-leaching operation is carried out at a temperaturebetween 425*F and 450*F.
 18. The process as defined in claim 1 whereinthe oxidation-leaching operation is carried out using an oxygen partialpressure of between about 100 and 200 pounds per square inch.
 19. Theprocess as defined in claim 1 wherein said neutralizing agent isselected from the group consisting of the hydroxides, oxides andcarbonates of calcium, strontium and barium.
 20. The process as definedin claim 1 wherein said neutralizing agent is calcium carbonate, whereinthe calcium in said calcium carbonate forms insoluble calcium sulfateand wherein said calcium sulfate is withdrawn from said autoclave aspart of the solids in said treated slurry.
 21. The process as defined inclaim 20 wherein said copper bearing sulfide ore also contains arsenic,wherein a first part of said calcium in said calcium carbonate formsinsoluble calcium sulfate, wherein said arsenic, solubilized to a metalsulfate in said oxidation-leaching operation, is precipitated with atleast a part of said calcium as insoluble iron and calcium arsenatesupon the addition of said calcium carbonate, and wherein said iron andcalcium arsenates are withdrawn from said autoclave as part of thesolids in said treated slurry.
 22. The process as defined in claim 1wherein the electrolytic copper produced is about 99.9 percent pure.